Methods for recovery of rare earth elements from coal

ABSTRACT

Methods of recovering rare earth elements, vanadium, cobalt, or lithium from coal are described. The coal is dissolved in a first solvent to dissolve organic material in the coal and create a slurry containing coal ash enriched with rare earth elements, vanadium, cobalt, or lithium. The enriched coal ash is separated from the first solvent. Residual organic material is removed from the coal ash. The rare earth elements, vanadium, cobalt, or lithium can then be recovered from the coal ash. The coal ash is mixed with an acid stream that dissolves the rare earth elements, thereby creating (i) a leachate containing the rare earth elements and (ii) leached ash. The leachate is heated to obtain acid vapor and an acid-soluble rare earth concentrate. The acid-soluble rare earth concentrate can be fed to a hydrometallurgical process to separate and purify the rare earth elements.

CROSS-REFERENCE TO RELATED APPLICATIONS

This application is a divisional of U.S. patent application Ser. No.15/617,589, filed on Jun. 8, 2017, which claims priority to U.S.Provisional Patent Application Ser. No. 62/347,194, filed on Jun. 8,2016, and to U.S. Provisional Patent Application Ser. No. 62/347,185,filed on Jun. 8, 2016. These applications are fully incorporated byreference in their entirety.

BACKGROUND

The present disclosure relates to the extraction and/or recovery of rareearth elements from coal ash, and systems and methods for doing so.

Rare earth elements (REEs) are a series of chemical elements found inthe Earth's crust. REEs generally include elements 21, 39, and 57-71 onthe periodic table. Due to their unique chemical properties, REEs havebecome critical in the development of many technologies spanning a rangeof applications including electronics, magnets, ceramics, computer andcommunication systems, transportation, health care, and nationaldefense. The demand and cost of REEs has grown significantly over recentyears, stimulating an emphasis on economically feasible approaches forREE recovery. Despite their criticality, however, there is currently notsignificant development of a domestic source of REEs.

Coal deposits in certain regions of the United States can beparticularly rich in rare earth elements, on the order of about 1,000parts-per-million (ppm). However, the combustion of coal in powerstations for energy generation yields coal ash as its product. Coal ashmay include fly ash and bottom ash from power generation stations, ashgenerated in a lower-temperature ashing process, and ash residual from acoal liquefaction process. The coal ash may contain REEs that areconcentrated above 300 ppm, and the REEs could then be economicallyrecovered. Processes for further concentrating and recovering REEs fromcoal ash would be desirable.

BRIEF SUMMARY

The present disclosure relates to acid digestion processes and systemsfor such acid digestion that can be used to recover and concentrate rareearth elements (REEs) from coal ash, as well as vanadium, lithium,and/or cobalt. Coal ash is preferred over raw coal, because the highorganic content of raw coal would consume acid used in the aciddigestion process, thereby making the process cost inefficient.Additionally, the concentration of REEs in raw coal or coal refuse istypically much lower than in coal ash.

In accordance with some illustrative embodiments disclosed herein,methods of recovering rare earth elements, vanadium, lithium, or cobaltfrom coal ash are disclosed. The methods include mixing coal ash with anacid stream such that rare earth elements present in the coal ash aredissolved in the acid stream, thereby creating (i) a leachate containingthe rare earth elements, vanadium, lithium, or cobalt, and (ii) leachedash, and heating the leachate to obtain acid vapor and an acid-solubleconcentrate containing the rare earth elements, vanadium, lithium, orcobalt. The acid-soluble rare earth concentrate may then be fed to ahydrometallurgical process to separate and purify the rare earthelements. The leached ash may be dried using heated air, which heatedair may also be used to heat the leachate. The heated air that was usedto dry the leached ash can include nitric acid with NO_(x) components.The leachate may be heated to dryness and then subsequently to a presettemperature. This preset temperature will selectively convert basemetals and/or rare earth target metals (i.e., REEs) in the leachate froma nitrate to an oxide. After this conversion, the solubility of thesolids in aqueous solutions will change, allowing for a later moreselective dissolution in the hydrometallurgical process to improveprocess economics. Additionally, conversion of metal nitrates to oxideswill release NO_(x) gases. Those NO_(x) gases may be recovered byoxidation in air and then absorption in a solvent such as cool water ordilute acid. The dried leached ash can be recovered. The acid stream canbe a nitric acid stream.

In accordance with some illustrative embodiments disclosed herein,systems for recovering rare earth elements, vanadium, lithium, or cobaltfrom coal ash are disclosed. The systems include a leaching reactor, anash dryer downstream of the leaching reactor, and a roaster downstreamof the leaching reactor. The roaster is cooperatively connected to boththe leaching reactor and the ash dryer. The system may further include asolid-liquid separation device downstream of the leaching reactor, withthe solid-liquid separation device feeding both the roaster and the ashdryer. A heat recovery exchanger may be present in the system betweenthe leaching reactor and the roaster. The heat recovery exchangerreceives the leachate containing rare earth elements from the leachingreactor and acid vapor recycled from the roaster. The system may alsoinclude a condenser downstream of the heat recovery exchanger. Thecondenser receives acid vapor from the heat recovery exchanger, andoutputs both a liquid acid stream and a residual vapor stream. Anabsorption column may be provided downstream of the condenser. Theabsorption column receives both the acid recovered from the condenserand the residual vapor stream. The acid in the vapor stream is captured,and then the acid is recycled to serve as the acid stream that is mixedwith the coal ash. The system may include a reactor heater upstream ofthe leaching reactor. The reactor heater provides preheated coal ash tothe leaching reactor.

In accordance with some illustrative embodiments disclosed herein, othermethods of recovering rare earth elements, vanadium, lithium, or cobaltfrom coal ash are disclosed. The methods include mixing coal ash with anacid stream in a leaching reactor such that rare earth elements presentin the coal ash are dissolved in the acid stream, thereby creating aleachate containing the rare earth elements and leached ash, and heatingthe leachate in a roaster to obtain acid vapor and an acid-soluble rareearth concentrate. The roaster process may include further concentratingthe leachate in a heat drum dryer using doctor knives to separate driedportions of the leachate, taking advantage of differential solubility ofthe dissolved metal salts to further separate and concentrate the rareearth elements. The heat drum dryer may be internally fired. Theacid-soluble rare earth concentrate may then be fed to ahydrometallurgical process to separate and purify the rare earthelements. The leached ash may be dried in an ash dryer using heated air,which heated air may also be used to heat the leachate in the roaster.The heated air can include nitric acid with NO_(x) components. Theleachate may be heated in the roaster to dryness and then subsequentlyto a preset temperature. This preset temperature will selectivelyconvert base metals and/or rare earth target metals (i.e., REEs) in theleachate from a nitrate to an oxide. After this conversion, thesolubility of the solids in aqueous solutions will change, allowing fora later more selective dissolution in the hydrometallurgical process toimprove process economics. Additionally, conversion of metal nitrates tooxides will release NO_(x) gases. Those NO_(x) gases may be recovered byoxidation in air and then absorption in a solvent such as cool water ordilute acid. The acid vapor from the roaster may be fed to a heatrecovery exchanger. At least a portion of the acid vapor fed through theheat recovery exchanger can be used to preheat the leachate before theleachate is heated in the roaster. At least a portion of the acid vaporfed through the heat recovery exchanger may be fed to a condenser thatrecovers acid for recycling. The acid recovered from the condenser andair containing residual oxide gases can be fed to a packed absorptioncolumn. The recovered acid can be cooled in the packed absorption columnand recirculated to absorb the oxide gases in the air and convert theoxide gases back to acid, thereby creating scrubbed air that issubstantially free of oxide gases. The recovered acid can be combinedwith a makeup acid stream. The combination of the recovered acid and themakeup acid stream can be recycled back to the leaching reactor. Nitricoxide (NO) gas generated in the leaching reactor may be oxidized to NO₂gas. The acid stream can be a nitric acid stream.

The present disclosure also relates, in accordance with someillustrative embodiments disclosed herein, to methods of recovering rareearth elements, vanadium, lithium, or cobalt from coal ash byliquefaction which are disclosed. By using coal ash from a coalliquefaction process, the REEs, vanadium, lithium, or cobalt can be moreeasily and economically leached, separated, concentrated, and recoveredby chemical or mechanical means (such as those described above).

The methods include dissolving coal containing rare earth elements,vanadium, lithium, or cobalt in a first solvent to dissolve organicmaterial in the coal and create a slurry containing coal ash enrichedwith REEs; separating the coal ash from the first solvent; removingresidual organic material from the coal ash; and recovering the rareearth elements from the coal ash. The first solvent may be bio-based,such as a bio-based hydrogen transfer solvent. The coal ash can beseparated from the first solvent by filtration, centrifugation, orsettling. The residual organic material can be removed from the coal ashby washing the coal ash with a second solvent, which may be differentfrom the first solvent. The residual organic material can also beremoved from the coal ash by burning the coal ash at a temperature ofabout 300° C. to about 600° C.

The methods may further include separating the coal ash into fractionscontaining the rare earth elements before recovering the rare earthelements from the coal ash. The coal ash can be separated into thefractions by density using a sink/float analysis. The coal ash canalternatively be separated into the fractions by particle size bysuccessively screening the coal ash. The coal ash could also beseparated into the fractions by particle size through the use of otherparticle sorting equipment, such as cyclones or air classifiers. Thecoal ash can also be separated into the fractions by chemical leaching.The chemical leaching can use a mineral base, an inorganic salt, or amineral acid. The mineral acid may be nitric acid. The chemical leachingcan be done by acid digestion. A calcination step can precede thechemical leaching.

In particular embodiments, the coal ash can be separated into thefractions by successive processes, such as by density using a sink/floatanalysis, then by particle size by successively screening the fractionsseparated by density, and then by chemical leaching of the resultingfractions.

The methods may further include purifying the rare earth elements,vanadium, lithium, or cobalt in a solvent extraction circuit to separateindividual elements from each other.

Also disclosed are methods of making a zeolite, comprising: adding azeolite seed to a leach solution containing silicon and aluminum; andheating the leach solution to obtain the zeolite. The leach solution canbe made by: mixing coal ash with a basic stream, thereby creating (i) aleach solution containing silicon and aluminum, and (ii) leached ash;and separating the leach solution from the leached ash. The leachsolution may be heated at a temperature of about 100° C. to about 200°C., and for a time of about 12 hours to about 96 hours.

These and other non-limiting characteristics are more particularlydescribed below.

BRIEF DESCRIPTION OF THE DRAWINGS

The following is a brief description of the drawings, which arepresented for the purposes of illustrating the exemplary embodimentsdisclosed herein and not for the purposes of limiting the same.

FIG. 1 is a schematic flow diagram of a closed-loop acid digestionprocess, namely a process of recovering rare earth elements from coalash.

FIG. 2 is a second schematic flow diagram of a closed-loop aciddigestion process for recovering rare earth elements from coal ash.

FIG. 3 is a bar graph showing the distribution of rare earth elementsafter treating at 200° C., in solution (left side) and in residualsolids (right side). The y-axis runs from 0% to 100% in increments of20%. The elements are, running from left to right, Sc, Y, La, Ce, Pr,Nd, Sm, Eu, Gd, Tb, Dy, Ho, Er, Tm, Yb, Lu, Fe, and Al. For all elementsexcept Sc, Fe, and Al, the amount in solution is higher than the amountin residual solids.

FIG. 4A and FIG. 4B are bar graphs showing the distribution of otherelements after treating at 200° C., in solution (left side) and inresidual solids (right side). The y-axis runs from 0% to 100% inincrements of 20%. In FIG. 4A, the elements are, running from left toright, Na, Li, Be, Mg, Si, K, Ca, Ti, V, Cr, Mn, Co, Ni, Cu, Zn, Ga, Ge,As, Se, Rb, and Sr. In FIG. 4B, the elements are, running from left toright, Zr, Nb, Mo, Ag, Cd, In, Sn, Sb, Te, Cs, Ba, Hf, Hg, Ta, W, TI,Pb, Bi, Th, and U.

FIG. 5 is a flow chart illustrating an exemplary method for recoveringrare earth metals from coal ash by liquefaction in accordance with someembodiments of the present disclosure.

FIG. 6 is a graph showing the REE concentrations (mg/L) of various REEsin different fractions. The y-axis runs from 0 mg/L to 800 mg/L inincrements of 100 mg/L. The elements are, running from left to right,Lu, Sc, Ce, Dy, Er, Eu, Gd, Ho, La, Nd, Pr, Sm, Tb, Tm, Y, Yb, HREE+Y,and LREE+Sc. With each element, from left to right, the fractions arethe feed coal, the low density and high density cuts, and four particlesize cuts (>850 microns, 600-850 microns, 355-600 microns, and <350microns).

FIG. 7 is a bar graph showing the ratio of heavy REE to light REE indifferent fractions. The y-axis runs from 0.0000 to 1.2000 in incrementsof 0.2000. From left to right along the x-axis, the fractions are thefeed coal, the low density and high density cuts, and four particle sizecuts (>850 microns, 600-850 microns, 355-600 microns, and <350 microns).

DETAILED DESCRIPTION

The present disclosure may be understood more readily by reference tothe following detailed description of desired embodiments. In thefollowing specification and the claims which follow, reference will bemade to a number of terms which shall be defined to have the followingmeanings.

Although specific terms are used in the following description for thesake of clarity, these terms are intended to refer only to theparticular structure of the embodiments selected for illustration in thedrawings, and are not intended to define or limit the scope of thedisclosure. In the drawings and the following description below, it isto be understood that like numeric designations refer to components oflike function.

The singular forms “a,” “an,” and “the” include plural referents unlessthe context clearly dictates otherwise.

The term “comprising” is used herein as requiring the presence of thenamed components/ingredients and allowing the presence of othercomponents/ingredients. The term “comprising” should be construed toalso include the term “consisting of”, which allows the presence of onlythe named components/ingredients, along with any impurities that mightresult from the manufacture of the named components/ingredients.

Numerical values should be understood to include numerical values whichare the same when reduced to the same number of significant figures andnumerical values which differ from the stated value by less than theexperimental error of conventional measurement technique of the typedescribed in the present application to determine the value.

All ranges disclosed herein are inclusive of the recited endpoint andindependently combinable (for example, the range of “from 2 grams to 10grams” is inclusive of the endpoints, 2 grams and 10 grams, and all theintermediate values). The endpoints of the ranges and any valuesdisclosed herein are not limited to the precise range or value; they aresufficiently imprecise to include values approximating these rangesand/or values.

The modifier “about” used in connection with a quantity is inclusive ofthe stated value and has the meaning dictated by the context. When usedin the context of a range, the modifier “about” should also beconsidered as disclosing the range defined by the absolute values of thetwo endpoints. For example, the range of “from about 2 to about 10” alsodiscloses the range “from 2 to 10.” The term “about” may refer to plusor minus 10% of the indicated number. For example, “about 10%” mayindicate a range of 9% to 11%, and “about 1” may mean from 0.9-1.1.

It should be noted that some of the terms used herein are relativeterms. For example, the terms “upstream” and “downstream” are relativeto the direction in which a fluid flows through various components, i.e.the flow fluids through an upstream component prior to flowing throughthe downstream component. It should be noted that in a loop, a firstcomponent can be described as being both upstream of and downstream of asecond component.

The present disclosure refers to pH. When referring to the pH of a givensolution as being “less than” a given value, this means the givensolution is more acidic than the given value, i.e. closer to a pH of 1.When referring to the pH of a given solution as being “greater than” agiven value, this means the given solution is more basic than the givenvalue, i.e. closer to a pH of 14.

The present disclosure refers to heavy rare earth elements, or HREE. Forpurposes of this disclosure, the “heavy” rare earth elements areyttrium, europium, gadolinium, terbium, dysprosium, holmium, erbium,thulium, ytterbium, and lutetium.

The present disclosure also refers to light rare earth elements, orLREE. For purposes of this disclosure, the “light” rare earth elementsare scandium, lanthanum, cerium, praseodymium, neodymium, promethium,and samarium.

Disclosed herein are systems and methods for recovering rare earthelements (REEs) from coal ash. The disclosed systems and methodsadvantageously provide an economical means of concentrating andrecovering REEs from coal ash. In addition, other elements such asvanadium, lithium, and cobalt can be recovered from the coal ash aswell. For purposes of this disclosure, discussion of REEs in generalshould also be considered as including vanadium, lithium, and cobalt.

With reference to FIG. 1, a diagrammatic representation of an aciddigestion process for recovering rare earth elements from coal ash isdepicted. Beginning in the upper left corner of FIG. 1, coal ash is fedinto the system through ash feed 102. It is specifically contemplatedthat the coal ash fed into the system can be high-temperature ash (i.e.,fly ash and/or bottom ash from power generation stations),low-temperature ash (i.e., ash generated in a lower-temperature ashingprocess), residual ash from a coal liquefaction process, or combinationsthereof. The coal ash is fed into the system at a suitable rate, such asabout 100 grams per minute (g/min) to about 500 g/min, including a rateof about 300 grams per minute, or at higher rates depending on thescaling of the system. For instance, it is contemplated that in afull-scale coal power plant, the coal ash feed could be on the order of1,000 tons per day. Coal ash particle diameters can range from severalsubmicrons up to several millimeters. The coal ash is generally composedprimarily of silica, alumina, and iron oxides, with remainder traceelements and unburned carbon deposits.

The coal ash is then mixed with an acid stream, which is a liquidstream, usually aqueous, containing an acid, i.e. a pH of less than 7.The acid is desirably nitric acid. However, it is contemplated thatother acids could be used, such as hydrochloric acid, sulfuric acid,hydrofluoric acid, and the like, and mixtures thereof. The acid streammay have a concentration of about 3M to about 8M. Desirably, thetemperature of the acid stream is somewhat elevated over roomtemperature, though this is not necessary. As illustrated here, the acidstream is a combination of recycled acid 168, make-up acid 108, andleachate recycle 141. The total acid stream can be fed at any desirablerate, for example from about 0.5 liters per minute (L/min) to about 5L/min, from any combination of these streams. The recycled acid 168 mayprovide an acid flow rate of about 0.5 L/min to about 1.5 L/min, and theleachate recycle 141 may provide an acid flow rate of about 0.5 L/min toabout 2.5 L/min.

The mixture of coal ash and acid can be fed to a reactor heater 112 topreheat the mixture. The reactor heater can preheat the mixture from atemperature of about 32° F. to about 68° F. (entering temperature) to atemperature of about 80° F. to about 212° F. (exiting temperature). Thereactor heater 112 then sends the preheated mixture to a leachingreactor 110. In this regard, the leaching reactor 110 is locateddownstream of the reactor heater 112.

The leaching reactor 110 is generally operated at a pressure of about 1atm and a temperature between about 20° C. and about 120° C. Within theleaching reactor 110, the coal ash is brought into intimate contact withthe acid stream and mixed therewith, such as with a static mixer or afluidized bed. Mixing of the coal ash with the acid stream causes REEs,V, Co, and Li present in the coal ash to be dissolved in the acidstream, thereby creating (i) a leachate containing the REEs, V, Co, andLi, and (ii) leached ash. The mixture may have a residence time withinthe leaching reactor of about 30 seconds to about eight hours. Ofcourse, this residence time may also be outside of this range, dependingon circumstances. Exiting the leaching reactor, the leachate and theleached ash are still mixed together. In particular embodiments, aneffective quantity of oxygen is introduced into the leaching reactor 110during mixing in order to oxidize NO_(x) gases. In this regard, nitrousoxide (NO) gas generated in the leaching reactor 110 can be oxidized toNO₂ gas.

Selectivity for the REEs, V, Co, and Li is higher than that for otherelements present in the coal ash (e.g., iron, aluminum, silicon),causing concentration of the REEs, V, Co, and Li in the acid stream overthe coal ash. The leaching reactor 110 generally causes an increase inthe surface area of the coal ash while removing some surfacecontaminants therefrom, which improves the pozzolanic activity of thecoal ash and makes it ideal for use in cements or other constructionmaterials.

The leached ash produced within the leaching reactor 110 can besubsequently filtered out from the leachate using a solid-liquidseparation device 140, such as a centrifuge, cyclone, settler, orfilter, or, for example, a ceramic microfilter. As shown in FIG. 1, thesolid-liquid separation device 140 is located downstream of the leachingreactor 110 and is fluidly connected thereto. In certain embodiments,the solid-liquid separation device 140 may be an air-scoured filterreceiving air from an air feed 104 that is fed by an air pump 170located upstream of the solid-liquid separation device 140. Two streamsexit the solid-liquid separation device 140, a leachate stream 142 and aleached ash stream 143. If desired, a portion of the leachate can berecycled upstream of the leaching reactor 110 or reactor heater 112, asindicated by reference numeral 141. This recycled leachate can be fedfrom the solid-liquid separation device 140 to the reactor heater 112 orthe leaching reactor 110 at a rate of, for example, about 1.5 L/min. Theleachate stream 142 is then sent to a roaster 130 at a rate of about 0.5L/min to about 5 L/min, depending on whether recycling is occurring.

The leached ash filtered out by the solid-liquid separation device 140subsequently travels to an ash dryer 120, where the leached ash is driedusing heated air. The heated air is produced by hot air feed 106. Theash dryer 120 is generally operated at a pressure of about 1 atm and atemperature of about 100° C. to about 300° C., for example 200° C. Theleached ash may have a residence time within the ash dryer of about twominutes to about one hour to be fully dried. Of course, the residencetime may also be outside of this range, depending on circumstances. Theash dryer 120 is important for economic recovery of the REEs. The hightemperatures experienced by the leached ash in the ash dryer 120generally cause boiling off and conversion of salts or esters (e.g.,nitrates) entrained in the leached ash, thereby allowing the salts oresters to be recovered in the system. This also prevents the dischargeof salts or esters from the leached ash wherever it is stored.

After the leached ash is dried in the ash dryer 120, it is separatedfrom the heated air by an ash separation process 122. Any suitableprocess for separating the dried ash from the heated air may beemployed. The dried ash and any other coal combustion byproducts maythen be recovered from the system (reference numeral 125). This mayoccur at a rate of about 95 g/m in to about 495 g/m in, depending on thescale of the system, and can be for example about 285 g/m in to about295 g/m in.

The heated air that is separated from the dried ash in the ashseparation process typically contains some nitric acid with NO_(x)components. As can be seen in FIG. 1, this heated air containing oxides(reference numeral 123) can be recovered from the ash dryer 120 and ashseparation process 122, and then fed to the roaster 130.

Returning to the leachate stream 142, a heat recovery exchanger 132 canbe present downstream of the leaching reactor 110 and upstream of theroaster 130. More specifically, the heat recovery exchanger 132 is shownin FIG. 1 as being located immediately downstream of the solid-liquidseparation device 140 and immediately upstream of the roaster 130. Theheat recovery exchanger 132 receives a portion of the leachate recoveredfrom the solid-liquid separation device 140 and preheats the leachateusing acid vapor 131 recovered from the roaster. The leachate thenenters the roaster 130.

The roaster 130 can be described as being located downstream of both theleaching reactor 110 and the ash dryer 120 and is cooperativelyconnected to both the leaching reactor 110 and the ash dryer 120. Theroaster 130 is also located immediately downstream of the heat recoveryexchanger 132 (i.e., downstream of the leaching reactor 110 with respectto the leachate). The roaster 130 is also located downstream of the ashdryer 120 with respect to the recovered heated air containing oxides123. That is, the roaster 130 receives recovered heated air from the ashdryer 120 and/or ash separation process 122, and receives the leachate142 from the leaching reactor 110, the solid-liquid separation device140, and/or the heat recovery exchanger 132. Additional acid may be fedinto the roaster via draw stream 151 if desired to help with heatrecovery. The roaster 130 is generally operated at a pressure of about 1atm and a temperature higher than that of the leaching reactor 110. Theoperating temperature of the roaster is high enough to convert amajority of the acid to the gas phase and to convert a majority ofcommon elements from a salt to an oxide, and in embodiments is about200° C. The roaster may take the form of, for example, an evaporatorfollowed by an internally-fired heat drum dryer.

The roaster 130 can be thought of as performing dual steps. First, theREEs (along with V, Co, and Li) are concentrated by drying the leachateto obtain metal salts. Next, the metal salts are crystallized by heatingto obtain metal oxides. In this regard, the leachate is generally heatedin the roaster 130 to dryness, and then subsequently to a presettemperature, which may be in the range of about 200° C. This presettemperature will selectively convert base metals in the leachate from anitrate to an oxide. In contrast, many of the target metals (i.e., REEs,V, Co, Li) remain in their nitrate form at this temperature. After theconversion from nitrate to oxide, the solubility of the solids inaqueous solutions will change, allowing for a later more selectivedissolution in the hydrometallurgical process to improve processeconomics. This permits removal of common elements (e.g. Al, Fe) fromthe REEs. Additionally, conversion of metal nitrates to oxides willrelease NO_(x) gases. These NO_(x) gases released by the metal oxidescan be subsequently recovered, as later described. The leachate may havea residence time within the roaster as needed to obtain the desired REEsin the desired form.

Heating of the leachate in the roaster 130 as described above results inan acid vapor 131 and an acid-soluble rare earth concentrate. 135. Theacid-soluble rare earth concentrate can typically be recovered from theroaster 130 at a rate of about 5 g/m in to about 15 g/m in, or higherdepending on the feed rate. The acid-soluble rare earth concentrate isparticularly suitable for feeding to a hydrometallurgical process toseparate and purify the REEs. Various hydrometallurgical processes canbe employed for separating and purifying the REEs, as will beappreciated by those skilled in the art. These processes may includeoxalate precipitation to purify rare earths, or more preferably solventextraction to purify and separate rare earths into distinct products.

During testing of the systems and methods disclosed herein, it wasobserved that bands of concentrated REEs with varying characteristicsformed during roasting of the leachate in the roaster 130. Accordingly,it is specifically contemplated that the roaster can include doctorknives that separate dried portions of the leachate from the REEs tofurther concentrate the REEs prior to the hydrometallurgical process. Asexplained previously, the roaster may take the form of, for example, anevaporator followed by an internally-fired heat drum dryer. In thisregard, it has been found that by finishing the drying of the rare earthconcentrate on the outside of a heated drum, the metal salts willprecipitate at different locations. This can then be scraped to obtainportions that are enriched with certain desirable REEs.

With continued reference to FIG. 1, the acid vapor 131 obtained fromheating the leachate in the roaster 130 is compressed and recycled.Recovery of the NO_(x) gases advantageously reduces the amount of nitricacid required by the system. In FIG. 1, the acid vapor is specificallyrecycled back through the heat recovery exchanger 132, upstream of theroaster 130. In this way, the acid vapor preheats the leachate that isfed to the roaster 130 (i.e., to preheat the leachate).

After passing through the heat recovery exchanger 132, the acid vaportravels to a condenser 150 for recovery of the acid for recycling. Asshown in FIG. 1, the condenser 150 is located downstream of the roaster130 and the heat recovery exchanger 132, and receives acid vaportherefrom. In particular embodiments, the acid vapor contains NO_(x)gases and nitric acid.

The condenser 150 is generally operated at a pressure of about 1 atm anda temperature of between about 40° C. and about 50° C. As explainedabove, the temperature at which the roaster 130 is operated is generallyabout 200° C. This large temperature differential between the roaster130 and the condenser 150 is important for improving acid recoverygreater than 90% by the condenser 150. In this regard, as the acid vaporpasses through the condenser 150, it is cooled, which is preferable forthe absorption of NO₂ back into the liquid phase. Recovery of the acidadvantageously reduces the amount of new acid required to be provided tothe system. Feasibility testing of the systems and methods disclosedherein evidenced recovery by the condenser 150 of greater than 90% ofthe acid used in the system.

When the acid stream used in the system is nitric acid, the NO gasgenerated in the leaching reactor 110, roaster 130, and ash dryer 120must be oxidized to NO₂ prior to being absorbed back into the acidstream. It has been discovered that this oxidation rate is improved athigh temperatures (e.g. 200° C.) and can occur during heating in theroaster 130 and drying in the ash dryer 120 with the introduction of aneffective quantity of oxygen therein.

Referring again to FIG. 1, the condenser 150 outputs a recovered acidstream 153 and a gas stream 152. The gas stream is generally aircontaining residual oxide gases (e.g., NO_(x) gases). The gas stream 152is fed to an absorption column 160, such as a packed absorption column.The packed absorption 160 column is located downstream of the condenser150. The absorption column 160 is generally operated at a pressure ofabout 1 atm and a temperature of between about 20° C. and about 30° C.As explained above, the temperature at which the condenser 150 isoperated is between about 40° C. and about 50° C.

Within the absorption column 160, the gas stream is contacted with asolvent, such as cool water or dilute acid, to recover the remainingacid. As illustrated here, the solvent is the recovered acid 153, whichcan also be recycled via line 163. This converts the oxide gases back toacid, thereby creating scrubbed air 173 that is substantially free ofoxide gases. As shown in FIG. 1, recirculation of the recovered acid canbe achieved using a recirculation pump 164, and cooling of the recoveredacid can be achieved using a column cooler 162. The scrubbed air is thenready to be discharged. Any residual oxide vapors can be recycled backto the absorption column 160, such as is shown in FIG. 1 via line 174,typically a flow rate of between about 15 g/min and about 20 g/min. Therecovered acid is then removed from the absorption column 160, typicallyat a flow rate of about 0.5 L/min to about 1.5 L/min. The recovered acid168 is generally combined with a makeup acid stream 108, as previouslydescribed, and recycled back as the starting acid that is mixed with thecoal ash 102. Most desirably, the system will only require less thanabout 2% makeup acid (i.e., the ratio of the makeup acid to recoveredacid).

FIG. 2 is a diagrammatic representation of another acid digestionprocess for recovering rare earth elements from coal ash. All of thecomponents of FIG. 1 are present, and additional components are added.

Beginning in the upper left corner of FIG. 2, coal ash is fed into thesystem through ash feed 202 to ball mill 210. It was found that millingthe ash achieved better leaching efficiencies. The ball mill reduces theparticle size of the ash from a range of 10 microns to 100 microns to arange of 1 micron to 40 microns.

The coal ash then travels downstream from ball mill 210 to caustic tank220 where it is mixed with a basic/caustic stream, which is a liquidstream, usually aqueous, containing a base, i.e. a pH of greater than 7.In this regard, the basic stream leaches silicon and aluminum (e.g. inthe form of silica and alumina) from the coal ash, thus increasing theconcentration of REEs and improving the leaching efficiency furtherdownstream. The base is desirably sodium hydroxide (NaOH). However, itis contemplated that other bases could be used, such as potassiumhydroxide, sodium bicarbonate, magnesium hydroxide, calcium hydroxide,and the like, and mixtures thereof. The basic stream may have aconcentration of 1% to about 10% of the base (w/w), though theconcentration can also be higher. Desirably, the temperature of thebasic stream is somewhat elevated over room temperature, though this isnot necessary. The basic stream can be fed at any desirable rate. Theash then travels from caustic tank 220 to caustic solid-liquid separator225, where the ash is separated from the basic stream. The basicsolution can be recycled to the tank 220 via line 227. The ash slurrytravels to leaching reactor 230 via line 228.

Alternatively, the basic solution filtered from the pretreatment of themilled coal ash can then be used to produce a zeolite material for useas an adsorbent, catalyst, catalyst support, or ion exchange media. Thisis done by heating the filtrate (i.e. basic solution) for a time afteradding a seed material, such as a zeolite crystal or a templatecompound. The heating temperature may be from about 100° C. to about200° C., and the heating time may be from about 12 hours to about 96hours. The resulting zeolite may contain aluminum and silicon, as wellas other elements such as potassium, sodium, oxygen, and hydrogen. Themolar ratio of silicon to aluminum in the resulting zeolite may be fromabout 4:1 to about 8:1. This is illustrated in FIG. 2, with line 294carrying the leach solution to processing tank 295 to produce thezeolite.

It is also noted that the functions of the ball mill 210 and the caustictank 220 could be combined together, with the ball milling taking placewithin a basic solution.

The leaching reactor 230 is downstream of the caustic tank 220. Theleaching reactor 230 of FIG. 2 serves the same function as leachingreactor 110 of FIG. 1. In the leaching reactor of FIG. 2, the coal ashis mixed with an acid stream. This mixing causes REEs present in thecoal ash to be dissolved in the acid stream, thereby creating (i) aleachate containing the REEs, V, Co, and Li, and (ii) leached ash.Again, an effective quantity of oxygen can be introduced into theleaching reactor 230 during mixing to oxidize NO_(x) gases. Exiting theleaching reactor, the leachate and the leached ash are still mixedtogether. Gases exit via line 231, which are sent to absorption column280.

The leached ash and the leachate are subsequently separated in asolid-liquid separation device 235, such as a centrifuge, cyclone,settler, or filter. The solid-liquid separation device 235 is locateddownstream of the leaching reactor 230 and is fluidly connected thereto.In certain embodiments, the solid-liquid separation device 140 may be adrum filter. Two streams exit the solid-liquid separation device 235, aleachate stream 239 and a leached ash stream 238. If desired, a portionof the leachate can be recycled back to the leaching reactor 230, asindicated by reference numeral 237.

The leached ash 238 is a slurry of ash and acid. This slurrysubsequently travels to an ash roaster/dryer 240, where the leached ashis dried. The dried ash and any other coal combustion byproducts maythen be recovered from the system (reference numeral 250).

The acid vapor, which may contain ash particles, travels from roaster240 through a particle separation device 244, such as an electrostaticprecipitator, to remove residual ash particles, which can also berecovered (reference numeral 250). The acid vapor then passes throughcondenser 246 to convert any acid from vapor into the liquid state, andthe liquid acid is subsequently sent to absorption column 280.

Returning to the leachate stream 239, the leachate stream 239 is sent toan evaporator 270 to remove additional acid from the leachate andfurther concentrate the REEs. The evaporator operates at a temperatureusually greater than 100° C. Acid vapor exiting the evaporator 270travels via line 271 to condenser 278, to convert the acid to the liquidstate, and the liquid acid then travels to absorption column 280 vialine 276.

The concentrated REE leachate, containing acid and REEs (and V, Co, andLi), then enters the roaster 260 via line 272. The roaster 260 can bedescribed as being located downstream of both the leaching reactor 230and the evaporator 270. The roaster 260 is generally operated at apressure of about 1 atm and a temperature of about 150° C. to about 200°C. The operating temperature of the roaster is high enough to convert amajority of the acid to the gas phase and to convert a majority ofcommon elements from a salt to an oxide. The roaster can be equipmentsuch as a spray dryer or rotary kiln.

The acid digestion in the leaching reactor 230 leads to formation ofnitrate salts with the general molecule formula M(NO₃)_(x), whereM(Al³⁺, Si²⁺, Sc³⁺, Eu³⁺ . . . ) is the cation extracted from fly ashand X is the valence: +1 (for Na, K, . . . ), +2 (Ca, Mg, Sr, . . . )and +3 (Ce, La, Cu, Fe, Al, Si, . . . ). In the roaster 260, the variousmetals in the leachate are dried into metal salts (i.e. nitrates). Thenitrate salts are then thermally decomposed to metal oxides, which canbe insoluble in water. In particular, iron and aluminum nitratesdecompose at temperatures of about 100° C. to about 200° C., while REEnitrates decompose at temperatures of about 300° C. to about 400° C.Sodium and calcium nitrates decompose at temperatures above about 400°C. As a result, iron and aluminum can be separated from the REEs inlater processing steps.

Heating of the leachate in the roaster 260 as described above results inan acid vapor and an acid-soluble rare earth concentrate 265 which canbe recovered and further processed in later downstream steps. The acidvapor obtained from heating the leachate in the roaster 260 travels tocondenser 275 to be converted to the liquid state, and the liquid acidthen travels to absorption column 280 via line 276.

The absorption column 260 receives a liquid acid stream 276 whichcontains acid recovered from roaster 240, roaster 260, and evaporator270 via condensers 246, 275, 278, respectively. The absorption column260 also receives a gas stream 231 from leaching reactor 230. Within theabsorption column 280, the gas stream is contacted with a solvent, suchas cool water or dilute acid. This converts the oxide gases back toacid, thereby creating an air stream that travels through a scrub 285 tocapture any residual NOx gases and results in scrubbed air 286, whichcan be discharged. The recovered acid is then removed from theabsorption column 160 and purified and/or combined with a makeup acidstream to obtain purified acid 291 which can be recycled back as thestarting acid that is mixed with the coal ash in leaching reactor 230.

Also disclosed herein are other methods for recovering rare earthelements (REEs) from coal. Ideally, the coal has been liquefied toobtain coal ash enriched with REEs. Other terms that can be used todescribe the coal ash include char or liquefaction residuals. Thedisclosed methods advantageously provide an economical means ofconcentrating and recovering REEs from coal by liquefaction.

FIG. 5 is a flowchart showing a general method 500 for recovering REEsfrom coal by liquefaction. The methods include receiving coal containingREEs 510. In particular embodiments, the coal is subjected to aliquefaction process to obtain coal ash enriched with REEs. By usingcoal ash from a liquefaction process, it has been found that REEs in thecoal ash can be more easily and economically leached, separated,concentrated, and recovered by chemical or mechanical means, such asthose disclosed above in FIG. 1 and FIG. 2.

The coal 510 is subsequently dissolved in a first solvent 520.Dissolution of the coal in the first solvent 520 dissolves organicmaterial in the coal and creates a slurry containing coal ash enrichedwith REEs. In certain embodiments, the first solvent may be bio-based,such as soybean oil. Bio-based solvents are oils derived from plants oranimals (rather than being derived from crude oil). It is to beunderstood that the first solvent can be any suitable solvent capable ofcausing dissolution of the coal. Desirably, the temperature of thecoal/first solvent mixture can be between about 300° C. to about 550°C., and the pressure can be from about 400 psi to about 1200 psi. Theresidence time can be between about 5 minutes to about 120 minutes.

After dissolution of the coal in the first solvent 520, the coal ashobtained therefrom is subsequently separated from the first solvent 530.Separation of the coal ash from the first solvent 530 can beaccomplished by any suitable means. In particular embodiments, the coalash is separated from the first solvent by a separation process selectedfrom the group consisting of filtration, centrifugation, or settling.The organic material remains in the first solvent, and the first solventcan be further processed to obtain coal liquid products such as fuels orchemical feedstocks.

Once the coal ash is separated from the first solvent 530, any residualorganic material in the coal ash is removed 540. Removing the residualorganic material from the coal ash 540 can be accomplished by anysuitable means. For example, in some embodiments, the residual organicmaterial is removed from the coal ash 540 by washing the coal ash with asecond solvent, such as tetrahydrofuran. The second solvent used to washthe coal ash is typically different from the first solvent used to causedissolution of the coal.

In other embodiments, the residual organic material is removed from thecoal ash 540 by burning the coal ash at a temperature sufficient tocause removal of the residual organic material, such as by burning thecoal ash at a temperature from about 300° C. to about 600° C. Othermethods for removal of the residual organic material from the coal ashmay be used, such as comminution, froth flotation, or gravimetricseparations.

After the residual organic material is removed from the coal ash 540,other processes can be used to concentrate/enrich the REEs, such asthose processes described in FIG. 1 and FIG. 2.

In certain embodiments, after removing the residual organic materialfrom the coal ash 540, the coal ash may be separated into fractions 550,some of which may preferentially contain REEs or preferentially containparticular REEs. Separating the coal ash into fractions 550 can beaccomplished by any suitable means. For example, in some embodiments,the coal ash is separated into fractions 550 by density using asink/float analysis. In other embodiments, the coal ash is separatedinto fractions 550 by particle size by successively screening the coalash, or using other particle sorting equipment, such as air classifiersor cyclones. In yet other embodiments, the coal ash is separated intofractions 550 by chemical leaching. The chemical leaching can use amineral base, such as sodium or potassium hydroxide, carbonates, orammonia. In other embodiments, the chemical leaching uses a mineral acidor an inorganic salt. The mineral acid may be any suitable acid capableof separating the coal ash into fractions. In yet other embodiments, thechemical leaching is performed by acid digestion. The acid digestion canbe performed by any acid, such as nitric acid, etc. The acid may have aconcentration of about 3M to about 8M. Desirably, the temperature of thecoal ash/acid is somewhat elevated over room temperature, though this isnot necessary. The temperature can be from about 20° C. to about 120° C.Such processes are described above. In yet further embodiments,separation of the coal ash into fractions can be accomplished using asalt solution such as ammonium sulfate to remove REEs from the ash.

In particular embodiments, the chemical leaching may be preceded by acalcination step. It is also to be understood that, in particularembodiments, the coal ash is separated into fractions 550 by successiveprocesses. That is, in certain embodiments, the coal ash is separatedinto fractions 550 by density using a sink/float analysis, then byparticle size by successively screening the fractions separated bydensity, and then by chemical leaching of the resulting fractions.

In some embodiments, the step of separating the coal ash into fractions550 is omitted, and a step of recovering the REEs from the coal ash 560immediately follows the step of removing the residual organic materialfrom the coal ash 540. In other embodiments, the step of separating thecoal ash into fractions 550 is employed prior to recovering the REEsfrom the coal ash 560. Recovery of the REEs from the coal ash 560 may beaccomplished by any suitable means, such as leaching, or any othersuitable chemical or mechanical means as will appreciated by thoseskilled in the art.

In certain embodiments, after recovering the REEs from the coal ash 560,the REEs are purified in a solvent extraction circuit to separateindividual REEs 570 from each other. Separation of the REEs isparticularly advantageous when the coal ash contains multiple differentREEs.

The following examples are provided to illustrate the systems andmethods of the present disclosure. The examples are merely illustrativeand are not intended to limit the disclosure to the materials,conditions, or process parameters set forth therein.

EXAMPLES Example 1

In feasibility studies using the method of FIG. 1, 10 grams of coal ashin 255 mL of dilute acid produced roughly 0.6 grams of roasted product.

Example 2

Key rare earth elements (REEs) include scandium (Sc), vanadium (V),neodymium (Nd), yttrium (Y), dysprosium (Dy), terbium (Tb), andpraseodymium (Pr). Scandium is primarily used as an alloying compoundwith aluminum to make high performance, lightweight alloys. Vanadium isused largely as an alloying compound to strengthen steel. but is alsouseful as a catalyst.

Tests were performed using a starting fly ash having the followingconcentrations:

Elements Concentration (ppm) REE + Y + Sc 545 ± 13.4 Sc 36 ± 1.4 V 279 ±12   Y 104 ± 5.3  Co 44 ± 2.5 Li ~166

Tests were performed at multiple nitric acid concentrations: 17%, 34%,51%, and 68% (w/w). The tests were performed on unmilled fly ash withparticle size of about 10 microns to about 100 microns, although onetest was done with milled fly ash having a particle size of about 1micron to about 40 microns. The results are shown in the followingtable. They indicate reduced leaching efficiency at higher acidconcentrations, which is likely due to passivation of the bulk aluminumand iron phases preventing further leaching. Aluminum and iron leachefficiency averaged 11.5% and 6.1%, respectively, in the 17% and 34%acid concentrations, compared to 3.4% and 2.4% at the higher acidconcentrations.

Starting Nitric Acid Concentration in PCC Fly Ash Leaches 34% Element17% 17% 17% 34% 51% 68% (milled) Sc 19.2% 20.8% 21.5% 21.5% N/A N/A55.3% Y 24.6% 26.7% 28.0% 28.0% 14.9% 13.0% 46.9% La 19.0% 19.3% 20.0%19.0% 9.9% 8.2% 35.4% Ce 21.0% 21.5% 21.7% 27.0% 11.9% 9.9% 34.0% Pr20.3% 21.7% 22.4% 22.9% 11.6% 10.0% 36.3% Nd 20.8% 22.6% 23.4% 23.9%12.3% 10.5% 39.5% Sm 22.5% 24.0% 25.0% 25.4% 13.7% 11.8% 40.5% Eu 22.7%24.5% 25.4% 26.4% 14.8% 12.7% 42.4% Gd 25.0% 27.2% 28.5% 28.8% 15.7%13.7% 45.2% Tb 23.3% 25.5% 26.9% 28.1% 15.4% 13.4% 44.3% Dy 24.1% 26.2%27.6% 28.6% 15.5% 13.0% 41.9% Ho 24.6% 26.8% 28.0% 28.6% 15.2% 13.3%41.8% Er 23.8% 26.2% 27.5% 27.8% 14.8% 12.6% 43.8% Tm 23.0% 25.2% 26.4%26.9% 14.4% 12.0% 42.2% Yb 21.2% 23.1% 24.7% 24.8% 12.9% 10.6% 36.3% Lu21.2% 22.6% 23.9% 24.3% 13.0% 10.2% 34.6%

Next, caustic pretreatment (tank 220 in FIG. 2) was tested. There weresix tests performed. The caustic solution used was sodium hydroxide, andthree different concentrations were tested (10%, 5%, and 1% NaOH) at twodifferent temperatures (20° C. and 90° C.). Each pretreatment was donewith a residence time of one hour, and unmilled fly ash was used as thestarting material. After pretreatment, leaching with 34% nitric acid wasperformed at 90° C. with a residence time of 30 minutes. After leaching,a sample was taken for analysis of scandium via Inductively CoupledPlasma Optical Emission Spectrometry (ICP-OES). An extra test 7 was doneusing milled fly ash. The conditions and results for each test are shownin the following table.

Caustic Acid Leach Temperature/ Temperature/ Concentration ReactionConcentration Reaction Scandium % Test of NaOH (w/w) time of nitric acidtime leached 1 10% 20° C./1 hour 34% 90° C./30 min 23.17% 2 10% 90° C./1hour 34% 90° C./30 min 54.27% 3 5% 20° C./1 hour 34% 90° C./30 min23.05% 4 5% 90° C./1 hour 34% 90° C./30 min 38.08% 5 1% 20° C./1 hour34% 90° C./30 min 23.40% 6 1% 90° C./1 hour 34% 90° C./30 min 21.77% 710% 90° C./1 hour 34% 90° C./30 min 88.21% (milled ash)

Next, tests were performed on leachate from fly ash. 100 grams of flyash was leached in nitric acid for 24 hours, then the slurry wasfiltered via 0.22-micron filter. The fraction of dissolved material was11.5% without any pretreatment of ash. The dissolved material obtainedfrom acid leaching was treated at a single temperature of 200° C. Thesolids and water leach solutions were also analyzed by ICP-MS aftertreatment at 200° C. FIG. 3, FIG. 4A, and FIG. 4B report the %distribution of each element in the two fractions.

FIG. 3 suggests that most of the REE nitrates produced by leaching offly ash in nitric acid solution are not decomposed to oxides after 200°C. heat treatment. However, most of the aluminum and iron decomposed tooxides, therefore they can be separated from the REEs. Around 70% ofscandium is converted to oxide at 200° C., and about 30% remains in thenitrate form and can be separated from the aluminum and iron.

FIG. 4A and FIG. 4B show the % distribution for the other elementsbetween the water leach and residual solids. Most of the titanium,vanadium, chromium, niobium, molybdenum, indium, tin, tungsten, andantimony nitrates decompose to insoluble oxides and therefore canseparated from REEs. Other elements such as manganese, gallium, and leaddecomposed partially to insoluble oxides. These results show thatroasting can be used to separate REEs from other elements.

Finally, the effect of pH on the recovery of REEs was tested. Extractiontests used leach solution from fly ash, which was dried using a heatinglap (about 150° C.). Then, dry material was leached in deionized water(250 mL), forming a loaded solution. The loaded solution was adjusted todifferent pHs as indicated below. The loaded solution was then extractedusing 15% CYANEX 572 in Solvent 467 diluent. CYANEX 572 is a phosphorusbased chelating extractant formulated for the extraction andpurification of rare earth elements. Solvent 467 is the trade name for akerosene range aliphatic diluent. However, in practice, any high boilingsolvent in which the extractant is soluble can be used, includingaromatic diluents.

For the calculation of percent recovery in the extractant (organicphase), a mass balance was performed using the results from ICP-MSanalysis done on the aqueous phase and using as starting material themass of each species in the solution loaded from leaching and roastingtests. Also, a negative mass in extractant was calculated for somespecies due to analytical error, so any negative mass calculated wasassumed to be zero. Furthermore, some results obtained from ICP-MSanalysis were below detection limits. Consequently, calculations wereperformed using the given detection limit. The following table shows thepercent recoveries of rare earths and other species in the extractantafter extraction of the solution at different pH. This table also showsthe purity or selectivity at different pHs.

% Recovery Species pH 1.03 pH 2.04 pH 2.51 pH 3.34 pH 4.01 pH 4.48 pH4.99 Sc 60.0% 90.00% 90.00% 50.00% 98.00% 60.00% 98.00% Y 68.4% 99.79%99.78% 99.78% 99.88% 99.98% 100.00% La 0.0% 0.00% 0.00% 0.00% 72.02%86.86% 98.28% Ce 0.0% 11.75% 7.43% 23.74% 95.23% 98.62% 99.90% Pr 0.0%2.39% 5.41% 44.90% 97.20% 99.25% 99.95% Nd 0.0% 14.09% 18.21% 61.51%98.12% 99.47% 99.97% Sm 0.0% 79.28% 82.87% 95.43% 99.70% 99.89% 100.00%Eu 0.0% 91.43% 92.91% 98.15% 99.81% 99.80% 99.99% Gd 0.0% 92.40% 94.07%97.68% 99.77% 99.93% 100.00% Tb 0.0% 98.09% 98.27% 99.08% 99.83% 99.69%99.98% Dy 31.2% 99.49% 99.57% 99.67% 99.89% 99.95% 100.00% Ho 51.4%99.73% 99.74% 99.70% 99.90% 99.76% 99.99% Er 74.4% 99.58% 99.60% 99.67%99.90% 99.91% 100.00% Tm 91.5% 99.83% 99.83% 99.15% 99.90% 99.32% 99.97%Yb 97.2% 99.95% 99.95% 99.85% 99.91% 99.88% 99.99% Lu 97.4% 99.77%99.77% 98.84% 99.88% 99.07% 99.95% Fe 89.6% 97.39% 97.39% 86.95% 99.48%89.56% 99.48% Al 7.1% 0.00% 0.00% 0.80% 98.80% 99.47% 99.97% Si 60.0%90.00% 90.00% 50.00% 98.00% 60.00% 98.00% REE + Y + Sc out of 24.4%50.88% 50.90% 60.98% 95.14% 95.93% 99.66% total REE + Y + Sc AvailableREE + Y + Sc out of 1.37% 4.18% 4.19% 7.15% 0.88% 1.04% 1.04% totalmeasured species (purity or selectivity) REE + Y + Sc out of 2.11%18.45% 18.76% 18.51% 0.97% 1.12% 1.17% total measured species excludingSilica (purity or selectivity)

In summary, ball milling and caustic pretreatment of the ash allowed forhigh recovery of REEs, with leaching efficiencies for scandium as highas 86% and near complete recovery of total REEs as a weighted average.Milling of the ash to a particle size of about 4 microns to about 5microns allowed these recoveries to be realized with a contact time ofabout one hour with 10% sodium hydroxide solution at 90° C., andleaching in 34% nitric acid for 30 minutes at 90° C.

A first recovery of the rare earth elements by thermal roasting of theloaded acid can oxidize the iron and aluminum between 100° C. and 200°C., generating an insoluble oxide material. In testing with actual leachsolutions, 90% of the REEs could be recovered from the roasted solidswith a water leach, while omitting over 90% of the iron and aluminum,and over 60% of the uranium and thorium. The water leach had aconcentration of 1.2% REEs, effectively leading to over a 20× increasein purity of the REEs over the fly ash feed.

Solvent extraction testing suggested that extraction for REEs was mostselective at a pH of 3.4 (i.e. between 3 and 4), where 61% of REEs wereextracted at over 7% purity (over 120× concentration over the feed flyash). pHs between 1 and 5 generally permitted good recovery. The primarycontaminants were sodium, aluminum, silica, calcium, and iron, butsodium, potassium, magnesium, and calcium were largely excluded from theextract. At pH 5, near quantitative REE recovery could be achieved (over99%), including less valuable lanthanum and cerium. These REE solutionscould then be separated with commercial operations such as furthersolvent extraction or ion exchange, or an emerging technology could beused such as electrowinning or electrophoresis.

Example 3

A bio-based coal liquefaction process according to FIG. 5 was applied tocoal. The process dissolves coal in a biosolvent, which prevents the ashfrom experiencing high temperature oxidizing environments. Afterdigestion of the coal, the resulting oil is centrifuged to remove ashand heavy carbon deposits, and the residual material was then analyzed.

The residual material was separated into two density cuts (high and low)by sink/float in deionized water. The residual material was alsoseparated into four particle size cuts to determine whether a simplemechanical separation could cause meaningful concentration of thesample. FIG. 6 shows the concentration of REE through the processing ofthe coal liquefaction ash. FIG. 7 indicates the ratio of heavy REE tolight REE in each cut, which also provides some information on changesin the ratio through the liquefaction process.

Referring to FIG. 6, it appears that the coal liquefaction processenriches REE over the feed coal. It is notable that the ash content ofthe coal liquefaction process samples used in this analysis was low; inthe range of 7% ash, where it is normally expected to be 30% ash orhigher. This could be due to preferential collection of larger particlesize ashes. Alternatively, the REE enrichment could be caused by lesserdissolution of coal components that bear REE elements, causingdifferential enrichment.

The coal liquefaction process also appears to enrich the ash with themore valuable heavy rare earth components, as indicated in FIG. 7.Although the liquefaction process does seem to enrich REE, simpledensity and particle size separations of the residual material did nothave a significant effect on REE enrichment.

The present disclosure has been described with reference to exemplaryembodiments. Obviously, modifications and alterations will occur toothers upon reading and understanding the preceding detaileddescription. It is intended that the present disclosure be construed asincluding all such modifications and alterations insofar as they comewithin the scope of the appended claims or the equivalents thereof.

The invention claimed is:
 1. A method of recovering rare earth elementsfrom coal, comprising: dissolving coal ash in a first solvent todissolve organic material in the coal and create a slurry containingcoal ash enriched with rare earth elements; separating the coal ash fromthe slurry; removing residual organic material from the separated coalash; and recovering the rare earth elements from the coal ash from whichresidual organic material has been removed.
 2. The method of claim 1,wherein the first solvent is a bio-based hydrogen transfer solvent. 3.The method of claim 1, wherein the first solvent is soybean oil.
 4. Themethod of claim 1, wherein the residual organic material is removed fromthe separated coal ash by washing the separated coal ash with a secondsolvent that is different from the first solvent.
 5. The method of claim4, wherein the second solvent is tetrahydrofuran.
 6. The method of claim1, wherein the residual organic material is removed from the separatedcoal ash by burning the coal ash at a temperature of about 300° C. toabout 600° C.; or wherein the residual organic material is removed fromthe separated coal ash by comminution, froth flotation, or gravimetricseparation.
 7. The method of claim 1, further comprising separating thecoal ash from which residual organic material has been removed intofractions containing the rare earth elements before recovering the rareearth elements from the coal ash from which residual organic materialhas been removed; wherein the fractions are separated by density using asink/float analysis, or by particle size by successively screening thecoal ash from which residual organic material has been removed, or byparticle size by air classifiers or cyclones, or by chemical leaching.8. The method of claim 7, wherein the chemical leaching uses a mineralbase, an inorganic salt, or a mineral acid, or wherein the chemicalleaching is performed by acid digestion.
 9. The method of claim 7,further comprising a calcination step before the chemical leaching. 10.The method of claim 1, further comprising purifying the rare earthelements in a solvent extraction circuit to separate individual elementsafter recovering the rare earth elements from the coal ash.
 11. Themethod of claim 1, wherein the rare earth elements are recovered fromthe coal ash from which residual organic material has been removed by:mixing the coal ash from which residual organic material has beenremoved with an acid stream such that the rare earth elements present inthe coal ash from which residual organic material has been removed aredissolved in the acid stream, thereby creating (i) a leachate containingthe rare earth elements, and (ii) leached ash; and heating the leachateto obtain acid vapor and a concentrate containing the rare earthelements.
 12. The method of claim 11, wherein the acid stream comprisesnitric acid, hydrochloric acid, sulfuric acid, or hydrofluoric acid. 13.The method of claim 11, further comprising treating the coal ash fromwhich residual organic material has been removed with a basic solutionprior to mixing the coal ash with the acid stream.
 14. The method ofclaim 13, further comprising milling the coal ash from which residualorganic material has been removed while the coal ash from which residualorganic material has been removed is treated with the basic solution.15. The method of claim 11, further comprising separating the leachateand the leached ash before heating the leachate.
 16. The method of claim15, further comprising drying the leached ash.
 17. The method of claim16, wherein residual acid in the leached ash is recovered as acid vaporthat is condensed and recycled to the acid stream.
 18. The method ofclaim 11, wherein the acid vapor obtained from heating the leachate iscondensed and recycled to the acid stream.
 19. The method of claim 11,wherein the leachate is heated to a temperature of about 150° C. toabout 200° C.
 20. The method of claim 11, further comprising feeding theconcentrate to a hydrometallurgical process to separate and purify therare earth elements.